thiosulfate leaching of gold – a review

Minerals Engineering 2001 14(2)

THIOSULFATE LEACHING OF GOLD – A REVIEW

MARK G AYLMORE AND DAVID M MUIR

CSIRO Division of Minerals, PO Box 90 Bentley WA 6982, Australia.

E-mail: [email protected]

(Received 26 May 2000 ;accepted 20 October 2000)

ABSTRACT

The ammoniacal thiosulfate leaching process for gold and silver extraction has been reviewed in terms of

leaching mechanism, thermodynamics, thiosulfate stability, and gold recovery options. The application to

different ore types and process options have also been discussed.

The thiosulfate leaching process it catalysed by copper and has several advantages over the conventional

cyanidation process. Thiosulfate leaching can be considered a non-toxic process, the gold dissolution

rates can be faster than conventional cyanidation and, due to the decreased interference of foreign

cations, high gold recoveries can be obtained from the thiosulfate leaching of complex and carbonaceous-

type ores. In addition, thiosulfate can be cheaper than cyanide.

The chemistry of the ammonia-thiosulfate - copper system is complicated due to the simultaneous

presence of complexing ligands such as ammonia and thiosulfate, the Cu(II)-Cu(I) redox couple and the

stability of thiosulfate in solution. However, by maintaining suitable concentrations of thiosulfate,

ammonia, copper and oxygen in the leach solution, and consequently, suitable Eh and pH conditions,

thiosulfate leaching can be made practical.

Generally the thiosulfate leaching conditions reported in the literature are severe with high reagent

consumption. Further investigations are required on leaching under low reagent concentrations over

extended periods where reagent consumption is low. For high grade ores or refractory sulfide ores where

some pretreatment processing is required to liberate gold, the in-situ generation of thiosulfate should be

investigated in more detail. This may lessen thiosulfate consumption and liberate more gold through the

oxidation of host sulfide minerals.

Cementation (or metal displacement), resins and to a limited extent, activated carbon can be used to

recovery gold from thiosulfate solutions. The use of resins to recover gold from solution appears to show

some promise, however more work is required to develop suitable elution and recovery methods, and

greater selectivity over copper.

While difficulties remain to be overcome, thiosulfate leaching has considerable potential as an effective

and less hazardous procedure for gold and silver extraction from auriferous ores.

Keywords

Gold ores, Precious metal ores, Hydrometallurgy, Leaching, Extractive metallurgy

mailto:[email protected]


2

INTRODUCTION

Considerable attention has been given to alternatives to the use of the conventional cyanidation/carbon in pulp

process to extract gold from gold ores. Interest in the use of non-cyanide methods for the dissolution of gold

arises from concerns regarding the toxicity of cyanide and the inability of cyanide solution to effectively leach

carbonaceous or complex ores. Sparrow and Woodcock (1995) have provided a summary review on non-cyanide

lixiviants for leaching of gold.

In addition to cyanide being hazardous, metal cyanide species derived from cyanide leaching (e.g. copper

cyanide) and other cyanide species end up in the tailings dam. This can lead to environmental problems where

cyanide pollutants reach the water table and cause destruction of animal life. This is a major problem where

residential areas are developing closer to the plant sites.

An alternative approach is to use thiosulfate to leach gold from ores. Thiosulfate is considered a non-toxic

alternative to cyanide and can leach gold faster than cyanide.

In the conventional cyanide process, the recovery of gold and silver from ores is hindered by a variety of metal

impurities like copper, arsenic, antimony, zinc and nickel, since these consume either cyanide or oxygen.

Leaching by thiosulfate decreases interference from these foreign cations. In fact the presence of copper in the

ore can be utilised directly in the leaching process.

High consumption of reagents and the lack of a cheap process for recovering gold have made thiosulfate leaching

uneconomical to date, in comparison with cyanide leaching. Consequently the process has not been widely used

on a commercial scale.

The following is a review of the ammoniacal thiosulfate leaching process. It covers the leaching mechanism,

thermodynamics, thiosulfate stability, and possible gold recovery processes. The application to different ore

types is also discussed. The discussion is largely restricted to ambient leaching conditions with limited reference

made to elevated temperature and pressure conditions where appropriate.

Development history

The recovery of precious metals using thiosulfate was first proposed early in the 1900s (White, 1900). In a

process known as the Von Patera process, gold and silver ores were first subjected to a chloridising roast and

then leached with thiosulfate. Silver-rich sulfide ores in South America were leached for many years prior to

World War II with thiosulfate after a chloridising roast (Flett et al., 1983). A similar treatment was also carried

out at the LaColorado Mine at Sonora in Mexico (Von Michaelis, 1987). However, it was not until the late

1970’s that an application to recover precious metals from copper-bearing metal sulfides concentrates and

pressure leach residues was developed employing ammonium thiosulfate and patented by Berezowsky and Kerry

(Hiskey and Atluri, 1988). During this period, studies were also carried out in the former Soviet Union (e.g. Ter-

Arakeyan et al., 1984). It was demonstrated that copper ions in solution could speed up the dissolution of gold.

Early research tended to concentrate on leaching at high temperatures and pressures to prevent copper sulfide and

sulfur layers from forming on gold particles thus preventing their leaching.

Berezowsky and Sefton (1979) revived interest in thiosulfate leaching by developing an atmospheric ammoniacal

thiosulfate leach process to recover gold and silver from residues of the ammonia oxidation leaching of sulfide

copper concentrates

More recent work has concentrated on understanding and improving the atmospheric ammoniacal thiosulfate

leach process. Extensive studies on the kinetics and mechanism of gold dissolution by thiosulfate have been

carried out by the Chinese and Japanese (e.g. Tozawa et al., 1981; Jiang et al., 1993).

Qian and Jiexue (1989) reported that a large plant had been built in Mexico using thiosulfate based on the Kerley

(1981, 1983) patents but could not be run successfully. Although these authors do not provide a reference to the

Mexican operation, they state that a subsequent patent by Perez (1987) claimed that the plant would have been

successful if run at pH 10.0-10.5 rather than pH 8 to avoid the problem observed with dissolution of iron from

milling media.

Heap leaching of carbonaceous preg-robbing ores using thiosulfate, has been the only industrial application in

recent years established by Newmont Gold Co. (Wan et al., 1994). Unlike gold cyanide, gold thiosulfate does not

adsorb on carbonaceous material and much higher gold recoveries can be achieved.


3

Conventional uses of thiosulfate

Its major industrial uses include the removal of excess chlorine in paper and textile bleaching, the preparation of

matches, the preservation of soap and as a chemical reagent.

Thiosulfate can be used as an antidote to cyanide and arsenic poisoning. It has been used to treat parasitic skin

disorders such as ringworm (Meyer, 1977). Ammonium thiosulfate has been used as a fertilizer for soils low in

sulfur for many decades. Environmentally it has distinct advantages over cyanide due to its low toxicity and

beneficial nature as a fertilizer. High concentrations of ammonium thiosulfate into water courses would, however

exasperate the common problem of excess nutrients and algae growth in rivers and lakes.

Although the price varies with location, sodium cyanide costs about US$2.00 per kilogram, whereas ammonium

thiosulfate costs about US$0.10 per kilogram. Therefore, providing reagent consumption is not high, thiosulfate

can be cheaper than cyanide.

GOLD DISSOLUTION IN THIOSULFATE LEACHING UNDER AMBIENT CONDITIONS

The chemistry of the ammonia-thiosulfate - copper system is complicated due to the simultaneous presence of

complexing ligands such as ammonia and thiosulfate, the Cu(II)-Cu(I) redox couple and the possibility of

oxidative decomposition reactions of thiosulfate involving the formation of tetrathionate and other additional

sulfur compounds (Umetsu and Tosawa, 1972; Kerley, 1981).

For clarity, the different components in the thiosulfate leaching system are discussed separately.

Thiosulfate leaching

In alkaline or near neutral solution of thiosulfate, gold dissolves slowly in the presence of a mild oxidant (White,

1905). The dissolution of the gold can be described as follows where oxygen is the oxidant and thiosulfate the

ligand.

4 Au + 8 S2O32-

+ O2 + 2H2O = 4[Au(S2O3)2]3-

+ 4 OH-

Two thiosulfate complexes of gold are known to form and these are Au(S2O3)- and Au(S2O3)2

3- with the latter

complex being the most stable (Johnson and Davis 1973).

Alkaline solutions must be used to prevent the decomposition of thiosulfate at low pH. This has the added

advantage of minimising the solubility of impurities, in particular iron compounds as discussed later. Once

formed, the thiosulfate complex is extremely stable. The stability constants for thiosulfate in comparison with

various other gold complexes are shown in Table 1.

TABLE 1 Stability constants for gold complexes

Gold species Stability Log K* Reference

Au(CN)2- 38.3 Smith and Martell, 1989

Au(SCN)2- 16.98 Smith and Martell, 1989

Au(SCN)4- 10 Smith and Martell, 1989

AuCl4- 25.6 Wang 1992

Au(NH3)2+ 26 Wang 1992

13 Hancock et al., 1974&

Au(S2O3)23-

26.5 IUPAC, 1993

28 Sullivan and Kohl, 1997

*Temperature 25C, &calculated from linear free energy relationship; Ionic strength = 1.0

Effect of ammonia

In the absence of ammonia, gold dissolution by thiosulfate is passivated by the build up of sulfur coatings as a

result of decomposition of thiosulfate on the gold surface (Pedraza et al., 1988; Jiang et al., 1993; Chen et al.,

1993). It is suggested that ammonia prevents gold passivation by being preferentially adsorbed on gold surfaces

over thiosulfate thus bringing gold into solution as an ammine complex (Jiang et al., 1993; Chen et al., 1996).


4

Ammonia reacts with gold ions forming gold ammonia complex ions that are then substituted by thiosulfate ions

(Chen et al., 1993).

The ammine complex is converted to thiosulfate complex as shown below

Au(NH3)2+ + 2S2O3

2- = Au(S2O3)2

3- + 2NH3

Although the dissolution of gold is thermodynamically feasible in ammoniacal solutions, kinetic experiments

have shown that gold is essentially not leached in ammonical solutions at room temperature (Meng and Han,

1993). Gold dissolution in ammonia solutions is observed only at temperatures above 80C.

The major role of ammonia in the thiosulfate system is to stabilise copper (II) as discussed below. However, the

presence of ammonia hinders the dissolution of iron oxides, silica, silicates and carbonates, the most common

gangue minerals found in gold bearing ores (Abbruzzese et al., 1995).

Effect of copper

The catalytic action of copper ions in promoting gold dissolution in thiosulfate solutions was first reported by

Tyurin and Kakowsky. (1960). Copper ions in solution can speed up the dissolution of gold by 18 to 20 times

(Ter-Arakeyan et al., 1984). In ammoniacal solutions and at temperatures below 60C, Cu(NH3)42+

is reported to

form (Tozawa et al., 1981). Gold dissolution can occur by using copper as the oxidant, rather than oxygen, as

follows

Au + Cu(NH3)42+

= Au(NH3)2+ + Cu(NH3)2

+

The redox equilibrium between the cuprous-cupric couple in ammoniacal solution and thiosulfate is represented

by the following reaction (Garrel et al., 1965; Abbruzzese et al., 1995; Wan, 1997)

2Cu(S2O3)35-

+ 8NH3 + 1/2O2 + H2O = 2Cu(NH3)42+

+ 2OH- + 6S2O3

2-

The role of copper(II) ions in the oxidation of metallic gold to aurous Au+ ion is shown in the following reaction

Au + 5S2O32-

+ Cu(NH3)42+

-> Au(S2O3) 23-

+ 4NH3 + Cu(S2O3)35-

The reported increase in dissolution of gold in copper thiosulfate solutions containing ammonia has been

attributed to the formation of copper (II) ammine complexes. This suggestion is supported by electrochemical

studies carried out by Chen et al. (1996). The copper catalytic process is described in more detail later.

In addition to the above processes, some thiosulfate degradation to tetrathionate occurs. The oxidation reaction,

which is promoted by copper (II) ion, is shown as follows

2Cu(NH3)42+

+ 8S2O32-

= 2Cu(S2O3)35-

+ S4O62-

+ 8NH3

Consequently, the concentration of copper (II) ion present in the leaching solution is an important factor in

thiosulfate stability and reagent management. The reduction of Cu (II) by thiosulfate ions is extremely rapid in a

pure aqueous solution but in the presence of ammonia the reduction reaction is slower and is dependent on the

ammonia concentration.

Effect of oxygen

Oxygen or some other oxidant is required to convert copper (I) to copper (II) for further gold leaching.

The most detailed work on the role of oxygen in ammoniacal thiosulfate systems containing copper (II) was

carried out by Byerley et al., (1973ab,1975). Depending upon the amount of oxygen dissolved in solution, rapid

oxidation of Cu(I) to Cu(II) occurs with some oxidation of thiosulfate to give sulfate and trithionate.

The oxidation of thiosulfate in aqueous solution by molecular oxygen, under ambient conditions of temperature

and pressure, is known to be extremely slow and only occurs when copper (II) ions and ammonia are present

(Naito et al., 1970).


5

In the absence of oxygen and under alkaline conditions, copper (II) ions in aqueous ammonia solution oxidise

thiosulfate ions initially to tetrathionate ions, the latter then undergoes a disproportionation reaction to yield

trithionate and thiosulfate ions. This reaction is catalysed by the presence of thiosulfate ions.

At low potentials where oxidants are deficient, in stagnant solutions, or in solutions containing high copper, the

decomposition of thiosulfate leads to the precipitation of black copper sulfides. Hence the precipitation of copper

sulfides is related to the availability of oxygen in the system. The limited solubility of oxygen in solutions and

the slow reduction at the gold surface makes the use of oxygen without the copper catalytic reaction very slow,

resulting in low gold dissolution.


6

THERMODYNAMICS OF GOLD DISSOLUTION

Thermodynamic calculations can help to clarify the observed experimental processes described above.

Zipperian et al (1988) constructed a series of Eh-pH diagrams to identify the predominant species that exist in

metal-ammonia-thiosulfate-water systems for high reagent concentrations (>1M). For low reagent

concentrations, Wang (1992) assumed the presence of AuNH3S2O3- ions in constructing Eh-pH diagrams

although the existence of AuNH3S2O3- has not been confirmed experimentally. However at low reagent

concentrations it could be expected since the stability constants of Au(NH3)2+ and Au(S2O3)2

3- complexes coould

be similar (see Table1).

The Eh-pH and speciation diagrams for ammoniacal thiosulfate and copper system, for high and low

concentrations of ammonia and thiosulfate are shown in the following Figures. The diagrams represent the

typical range of leaching conditions used. Construction of Eh-pH and speciation diagrams was carried out using

Outokumpu HSC Chemistry for Windows (Roine, 1994) and MINEQL+ software (Schecher and McAvoy,

1998). Thermodynamic data was obtained from Smith et al (1998) and is listed in Table 2.

TABLE 2 Free energies of formation (kJ/mol) for copper ammonical thiosulfate species

Species G298

(kJ/mol)

Species G298

(kJ/mol)

Species G298

(kJ/mol)

S 0 Ag 0 Cu 0

S5O62-

(a) -956.0 AgO 10.9 CuO -127.194

S3O62-

(a) -958.0 AgOH -92.0 Cu2O -146.356

S2O62-

(a) -966.0 Ag2O3 87.0 Cu(OH)2 -356.895

SO32-

(a) -486.5 Ag2O -10.8 CuS -48.953

HS2O3-(a) -541.8 Ag2S -40.5 Cu2S -86.1904

H2S(a) -27.3 Ag2+

(a) 268.2 Cu2+

(a) 64.978

HS-(a) 12.6 Ag

+(a) 77.2 Cu

+(a) 50.208

H2SO3(a) -537.9 AgO-(a) -23.0 Cu(S2O3)3

5-(a) -1624.65

H2S2O3(a) -543.5 Ag(S2O3)35-

(a) -1598.3 Cu(S2O3)23-

(a) -1084.07

H2S2O4(a) -436.3 Ag(S2O3)23-

(a) -1058.6 Cu(S2O3) -(a) -540.991

HSO3-(a) -527.7 Ag(S2O3)

-(a) -506.3 Cu(NH3)

+(a) -10.293

HS2O4-(a) -434.2 Ag(NH3)

+(a) 31.8 Cu(NH3)

2+(a) 14.477

S2-

(a) 91.9 Ag(NH3)2+(a) -17.5 Cu(NH3)2

2+(a) -32.259

S22-

(a) 79.7 Cu(NH3)32+

(a) -73.212

S32-

(a) 73.8 Au 0.0 Cu(NH3)42+

(a) -112.968

S42-

(a) 69.4 AuO2 200.8 CuO22-

(a) -181.167

S52-

(a) 66.1 Au(OH)3 -290.0 HCuO2-(a) -256.981

S2O32-

(a) -532.2 Au2O3 163.2

S2O42-

(a) -600.6 Au+(a) 163.2

S2O82-

(a) -1115.0 Au3+

(a) 433.5

S4O62-

(a) -1040.4 H3AuO3(a) -258.6

Au(S2O3)23-

(a) -1050.2

NH3(a) -26.7 Au(NH3)43+

(a) 64.4

NH4+(a) -79.5 Au(NH3)2

+(a) -41.4 or 35.7

AuO33-

(a) -24.3

HAuO32-

(a) -115.5

H2AuO32-

(a) -191.6

Increase in dissolution of gold in copper thiosulfate solutions containing ammonia is attributed to the

stabilisation of the oxidant Cu(II) by ammonia by forming copper (II) ammine complexes. Therefore the

conditions need to be chosen where both copper (II) ammine complex and gold dissolution occur.

It is clear in both Eh-pH conditions illustrated (Figures 1 and 2) that high pH values should be avoided since

copper will be removed from the leach solution as oxides. In addition, potential influences the equilibrium

between copper (I) and copper (II) ions. At high potentials ammonia copper (II) complexes exist, whereas at low

potentials copper (I) thiosulfate complexes exist.


7

Decreasing the ammonia, thiosulfate and copper (II) concentrations significantly narrows the region of stability

of Cu(NH3)42+

and Cu(S2O3)35-

, and expands the stability region of CuO, Cu2O and Cu2S.

As the concentrations of thiosulfate, copper and ammonia change the stability regions of gold species change

slightly. It should be noted that the pH values change significantly, with change in the concentration of either

thiosulfate or ammonia. If the potentials in a copper - ammoniacal thiosulfate system are too low, gold remains

undissolved over the whole pH range from 0-14. In addition, copper in solution will precipitate out as a sulfide.


8

(a)

141210864

2.0

1.5

-0.5

-1.0

-1.5

-2.0

Eh (Volts)

pH

Au(NH3)4+3 (aq)

Au

0.0

1.0

0.5 Au(S2O3)2-3 (aq) {Au(NH3)2

+(aq)}

(b)

Fig 1 Eh-pH diagram at high reagent concentrations for (a) Cu-NH3-S2O32-

system and (b) Au-NH3-S2O32-

system (conditions: 5x10-4

M Au; 1MS2O32-

;1M NH3/NH4+, 0.05M Cu

2+). The dotted line marks the

stability region of Au(NH3)2+ where the stability constant log K= 26.0 (see Table II). The shaded region

is the predominance area for Au(S2O3)23-

where stability constants are based on linear free energy

calculations (Hancock et al., 1974).

141210864

2.0

1.5

1.0

0.5

0.0

-0.5

-1.0

-1.5

-2.0

Eh (Volts)

pH

Cu

CuS

Cu2S

CuO

CuO

Cu2OCu(S2O3)3

-5 (aq)

Cu(NH3)4+2 (aq)

Cu+2 (aq)


9

(a)

1286

pH

AuO2Au(NH3)4+3 (aq)

4

2.0

-0.5

-1.0

-1.5

-2.0

Eh (Volts)

14

1.5

0.0

10

Au

Au(S2O3)2-3 (aq) {Au(NH3)2

+(aq)}

1.0

0.5

(b)

Fig 2 Eh-pH diagram at low reagent concentrations for (a) Cu-NH3-S2O32-

system and (b) Au-NH3-S2O32-

system

(conditions: 5x10-4

M Au; 0.1M S2O32-

; 0.1M NH3/NH4+, 5x10

-4M Cu

2+). The dotted line marks the

stability region of Au(NH3)2+ where the stability constant log K= 26.0 (see Table II). The shaded region

is the predominance area for Au(S2O3)23-

where stability constants are based on linear free energy

calculations (Hancock et al., 1974).

141210864

2.0

1.5

1.0

0.5

0.0

-0.5

-1.0

-1.5

-2.0

Eh (Volts)

pH

Cu

CuS

Cu2S

CuO

CuO

Cu2O

Cu(S2O3)3-5 (aq)

Cu(NH3)4+2 (aq)

Cu+2 (aq)


10

Eh-pH diagrams of Au-NH3 - S systems (Figures 1b and 2b) and species distribution diagrams (Figure 3) in

aqueous solution indicate that under certain conditions gold could be present in solution as [Au(NH3)2]+ rather

than as [Au(S2O3)2]3-

(Zipperian et al., 1988). The gold (I) thiosulfate complex is the most stable species in the

leaching system up to pH 8.5 or below 0.01M NH4+. Above this pH, when NH4

+ converts to NH3 and when the

ammonia concentration exceeds 0.1M, the predominate gold compound is gold (I) diammine complex (Zipperian

et al., 1988) (designated (a) in Figure 3).

0.00E+00

1.00E-05

2.00E-05

3.00E-05

4.00E-05

5.00E-05

-3 -2 -1 0 1 2 3 4 5 6

Log [NH3] (M)

Co

nc

en

tra

tio

n (

M)

Au(NH3)2+

Au(S2O3)23-

(a) (b)

Fig 3 Distribution of gold species at different ammonia concentrations (0.1M S2O32-

, 5x10-5

M Au, pH 9.5,

Eh=0.250V) with stability constant for Au(NH3)2+ at (a) Log K = 28 and (b) log K = 13.

However measurements at pH >9 have shown that gold rest potentials change with thiosulfate concentration

rather than with ammonia concentration (Wan, 1997). This suggests that the predominant gold species is

Au(S2O3)23-

rather than Au(NH3)2+. The variation between thermodynamics and electrochemical studies may be

attributed to the high activation energy for Au(NH3)2+ formation (Meng and Han, 1993). In addition, predicted

stability constants based on linear free energy relationship between silver (I) and gold (I) (Hancock et al., 1974)

derive a lower stability constant for Au(NH3)2+ than that for Au(S2O3)2

3- . Based on these calculations a much

higher concentration of ammonia is required to stabilised Au(NH3)2+ (designated (b) in Figure 3) and Au(S2O3)2

3-

exists over the whole pH range under conditions examined in Figures 1b and 2b (shaded area).

Eh-pH diagrams can be used to show the predominant species under different potentials and pH conditions.

However, speciation diagrams are required to characterise the distribution of various Cu-NH3-S2O32-

species co-

existing in solution.

The copper species distributions for high and low reagent concentrations at fixed Eh are shown in Figures 4a and

b respectively. It can be seen that to leach gold under low reagent conditions only a narrow pH region around pH

9.5 to 10.0 exists where copper ammonia complex is stable without the precipitation of copper (II) oxide,

tenorite. At high reagent concentrations a broader pH range is available.

Increasing the potential (e.g. with O2) but keeping all reagent concentrations constant results in the reduction in

Cu(S2O3)35-

and an increase in Cu(NH3)42+

in the solution (Figures 5a and b). At low NH3-S2O32-

-Cu2+

concentrations in solution, other copper ammonia and thiosulfate complexes are stable which can have a

significant effect on thiosulfate stability and leaching properties. Byerley et al., (1973) reported the role of the


11

triammine copper (II) complex rather than the tetraammine copper (II) complex in the oxidation of thiosulfate to

tetrathionate. In addition, only a small Eh range is available for maintaining copper ammonia complex in

solution.

Increasing total ammonia concentration but keeping the other reagent concentrations, pH and Eh constant,

increases the stability region of the Cu(NH3)42+

complex (6a and b), whereas increasing the thiosulfate

concentration ions increases the stability region of the Cu(S2O3)35-

complex (Figures 7a and b).

Under the conditions used, copper concentration has a much more pronounced effect on the stability regions of

copper species where Cu(S2O3)35-

, becomes more stable over the Cu(NH3)42+

complex for both high and low

NH3-S2O32-

-Cu2+

concentrations in solution (Figures 8a and b). In addition, precipitation of tenorite occurs with

increased copper concentration in solution. It would appear that a higher ammonia to thiosulfate concentration

would be required to achieve a higher Cu(NH3)42+

concentration in solution over Cu(S2O3)35-

(see Figure 6).

However, high Cu(NH3)42+

concentrations in solution will also result in higher losses of thiosulfate through its

conversion to tetrathionate.

It has been reported that significantly increasing the copper concentration can result in the precipitation of the

mixed copper(II)-copper(I) species Na[Cu(NH3)4. Cu(S2O3)2]. The addition of copper sulfate can result in the

precipitation of copper (I) thiosulfate and oxidation of thiosulfate to tetrathionate (Flett et al., 1983).

No inhibition of gold dissolution in thiosulfate solutions by excessive ammonia (0-2M) occurs in the absence of

copper (Chen et al., 1996). This precludes the widely held assumption that hydroxide ions in high concentration

that are produced by excessive ammonia, inhibit gold dissolution.

Regeneration of the copper (II) to sustain the catalytic reaction is an important feature of the leaching reaction

and so the concentration ratio of ammonia to thiosulfate has to be maintained so that copper can easily transfer

between the Cu(II) and Cu(I) states. For effective gold dissolution at high reagent concentrations, Zipperian et

al., (1988) determined that an oxidation potential of 150-200mV with reference to standard hydrogen electrode

(SHE) is required at pH 10. Eh-pH diagrams constructed here suggest a higher oxidation potential of 250mV is

required. At lower reagent concentrations the optimum Eh appears the same, but the pH range is smaller.


12

Fig 4a Distribution of copper species at different pH conditions for high reagent concentrations (1M S2O32-

, 1M

NH3, 0.05M Cu, Eh=0.250V).

Fig 4b Distribution of copper species at different pH conditions for low reagent concentrations (0.1M S2O32-

,

0.1M NH3, 5x10-4

M Cu, Eh=0.250V).

0.00E+00

1.00E-02

2.00E-02

3.00E-02

4.00E-02

5.00E-02

0 2 4 6 8 10 12 14

pH

Co

nc

en

tra

tio

n (

M)

Cu(NH3)42+

TenoriteCu(S2O3)35-

Cu(S2O3)23-

Cu(S2O3)-

Cu(OH)4-

0.00E+00

1.00E-04

2.00E-04

3.00E-04

4.00E-04

5.00E-04

0 2 4 6 8 10 12 14

pH

Co

nc

en

tra

tio

n (

M)

Cu(S2O3)35-

Cu(S2O3)23-

Cu(S2O3)-

Cu(NH3)42+

Tenorite

Cu(NH3)32+

Cu(OH)4-

Cu(OH)3-


13

0.00E+00

1.00E-02

2.00E-02

3.00E-02

4.00E-02

5.00E-02

0 0.1 0.2 0.3 0.4 0.5 0.6

Eh (V)

Co

nc

en

tra

tio

n (

M)

Cu(NH3)42+

Cu(S2O3)35-

Cu(S2O3)23-

Cu(NH3)32+

Fig 5a Distribution of copper species at different Eh conditions for high reagent concentrations (1M S2O32-

, 1M

NH3, 0.05M Cu, pH=10.0).

0.00E+00

1.00E-04

2.00E-04

3.00E-04

4.00E-04

5.00E-04

0 0.1 0.2 0.3 0.4 0.5 0.6

Eh (V)

Co

nc

en

tra

tio

n (

M)

Cu(NH3)42+

Cu(S2O3)35-

Cu(S2O3)23-

Cu(S2O3)- Cu(NH3)3

2+

Tenorite

Fig 5b Distribution of copper species at different Eh conditions for low reagent concentrations (0.1M S2O32-

,

0.1M NH3, 5x10-4

M Cu, pH 10.0).


14

0.00E+00

1.00E-02

2.00E-02

3.00E-02

4.00E-02

5.00E-02

0 1 2 3 4 5

Total NH3 conc (M)

Co

nc

en

tra

tio

n (

M)

Cu(NH3)42+

Cu(S2O3)35-

Cu(S2O3)23-

Cu(NH3)32+

Fig 6a Distribution of copper species at different NH3 conditions for high reagent concentrations (1M S2O32-

,

0.05M Cu, pH 10.0, Eh=0.250V). At pH 10.0 less than half the total ammonia is present as ammonium.

0.00E+00

1.00E-04

2.00E-04

3.00E-04

4.00E-04

5.00E-04

0 0.2 0.4 0.6 0.8 1

Total NH3 conc (M)

Co

nc

en

tra

tio

n (

M)

Cu(NH3)32+

Cu(NH3)42+

Cu(S2O3)35-

Cu(S2O3)23-

Cu(S2O3)-

Cu(NH3)2+

Fig 6b Distribution of copper species at different NH3 conditions for low reagent concentrations (0.1M S2O32-

,

5x10-4

M Cu, pH 10.0, Eh=0.250V). At pH 10.0 less than half the total ammonia is present as

ammonium.


15

0.00E+00

1.00E-02

2.00E-02

3.00E-02

4.00E-02

5.00E-02

0 1 2 3 4 5

S2O32-

conc (M)

Co

nc

en

tra

tio

n (

M)

Cu(S2O3)35-Cu(NH3)4

2+

Cu(S2O3)23-

Cu(NH3)32+

Fig 7a Distribution of copper species at different S2O32-

conditions for high reagent concentrations (1M NH3,

0.05M Cu, pH=10.0, Eh=0.250V).

0.00E+00

1.00E-04

2.00E-04

3.00E-04

4.00E-04

5.00E-04

0 0.2 0.4 0.6 0.8 1

S2O32-

conc (M)

Co

nc

en

tra

tio

n (

M)

Cu(S2O3)35-

Cu(S2O3)23-

Cu(S2O3)-

Cu(NH3)42+

Cu(NH3)32+

Tenorite

Fig 7b Distribution of copper species at different S2O32-

conditions for low reagent concentrations (0.1M NH3,

5x10-4

M Cu, pH 10.0, Eh=0.250V).


16

0.00

0.10

0.20

0.30

0.40

0.50

0.60

0.70

0.80

0 0.2 0.4 0.6 0.8 1

Cu2+

Conc (M)

Co

nc

en

tra

tio

n (

M)

Cu(S2O3)35-

Cu(S2O3)23-

Cu(NH3)42+

Tenorite

Cu(NH3)22+

Fig 8a Distribution of copper species at different Cu2+

conditions for high reagent concentrations (1M S2O32-

,

1M NH3, pH 10.0, Eh=0.250V).

0.0000

0.0001

0.0002

0.0003

0.0004

0.0005

0.0006

0.0007

0.0008

0.0009

0.0010

0 0.002 0.004 0.006 0.008 0.01

Cu2+

Conc (M)

Co

nc

en

tra

tio

n (

M)

Cu(S2O3)35-

Cu(S2O3)23-

Cu(NH3)42+

Tenorite

Cu(S2O3)2-

Cu(NH3)32+

Fig 8b Distribution of copper species at different Cu2+

conditions for low reagent concentrations (0.1M S2O32-

,

0.1M NH3, pH 10.0, Eh=0.250V).


17

THE ELECTROCHEMICAL-CATALYTIC MECHANISM OF GOLD LEACHING

The electrochemical-catalytic mechanism that is proposed to identify the leaching of gold with ammoniacal

thiosulfate is illustrated in Figure 9.

Fig 9 The model of electrochemical – catalytic mechanism of ammoniacal thiosulfate leaching of gold.

Thiosulfate ions react with Au+ ions on the anodic surface of gold and enter the solution to form Au(S2O3)2

3-. The

Cu(NH3)42+

present in solution acquires electrons on the cathodic portion of the gold surface and is directly

reduced to Cu(NH3)2+. In the presence of S2O3

2-, Cu(NH3)2

+ converts to Cu(S2O3)3

5- ions. The Cu(S2O3)3

5- species

in solution is then oxidised into Cu(NH3)42+

with oxygen (Wan, 1997). Likewise, the Cu(NH3)2+ species, if

present in solution, is oxidised into Cu(NH3)42+

with oxygen (Jiang et al., 1993). The predominant cathodic

reaction will depend upon the relative concentrations of the species in solution.

The mechanism is further complicated by the oxidation of some thiosulfate to tetrathionate. In the absence of

oxygen and pH>10, copper (II) ions in aqueous ammonia solution oxidise thiosulfate ions to tetrathionate ions.

Over a extended period of time, tetrathionate can further disproportionate to yield trithionate and thiosulfate ions.

The suggested mechanism from kinetic studies of the reaction indicates that substitution of thiosulfate ion into

the co-ordination sphere of ammine copper (II) complex occurs. At pH >10 , the copper(II) ammine species are

reported as existing in equilibrium with significant concentrations of hydroxo-species. Therefore, in addition to

tetra-ammine, some tri-ammine copper (II) species co-exist. An electron-transfer from the thiosulfate to the

copper (II) ion, occurring in the intermediate tri-ammine copper(II)-thiosulfate complex, gives rise to copper (I)

and S2O32-

ions, which in turn dimerise to tetrathionate ions (Byerley et al., 1973).

Gold surface

Anodic area

Au=Au+ + e

Au+ + 2S2O3

2- = Au(S2O3)2

3-

Au

Cathodic area

Cu(NH3)42+

+ e = Cu(NH3)2+ + 2NH3

Cu(NH3)2+ + 3S2O3

2- = Cu(S2O3)3

5- + 2NH3

2Cu(NH3)42+

+ 8S2O32-

= 2Cu(S2O3)35-

+ S4O62-

+ 8NH3

S2O32-

Au(S2O3)23-

Cu(NH3)42 +

S2O3

2-

Cu(NH3)2+ + 2NH3

e

+ O2 OH

-

Cu(S2O3)35-

+ NH3

+ S2O32-


18

The formation of tetrathionate from thiosulfate ions using Cu(NH3)42+

or Cu(NH3)32+

as oxidants can be

demonstrated by examining the relevant redox potentials for the systems. The calculated redox potentials from

G and stability constants are shown in Table 3.

TABLE 3 Redox potentials for copper ammonia and thiosulfate complexes

Couple Redox potential (V)

Cu(NH3)42+

/ Cu(S2O3)35-

0.22

Cu(NH3)42+

/ Cu(S2O3)23-

0.14

Cu(NH3)32+

/ Cu(S2O3)35-

0.36

Cu(NH3)32+

/ Cu(S2O3)23-

0.27

S2O32/S4O6

2- 0.12

It can be seen that redox potentials both for Cu(NH3)42+

/ Cu(S2O3)35-

and Cu(NH3)42+

/ Cu(S2O3)25-

are greater

than that for S2O32/S4O6

2 .

In the presence of oxygen, the Eh rises and rapid oxidation of Cu(I) to Cu(II) occurs with further oxidation of

thiosulfate to sulfate and trithionate, depending upon the amount of oxygen dissolved in solution. Detailed

kinetic studies undertaken by Byerley et al, (1973) suggested a mechanism in which O2 and thiosulfate become

associated with ammine-thiosulfato-copper (II) species and form sulfate and trithionate ions in solution. In this

role O2 assists in electron transfer between S2O32-

and Cu(II).

Clearly only sufficient O2 necessary to oxidise Cu(I) to Cu(II) is required to prevent the significant loss of

thiosulfate from solution.

Regeneration of the copper (II) to sustain the catalytic reaction is an important feature of the leaching reaction

and so the concentration ratio of ammonia to thiosulfate has to be maintained.

In principal, ammonia and cupric ions are not consumed and the reagent replenishment is limited to thiosulfate.

Loss of some ammonia from the leaching solution is unavoidable but can be reduced providing the pH and

aeration are properly maintained.

Tao et al (1993) and Michel and Frenay (1996) have evaluated by electrochemical means, a similar mechanism

of gold oxidation and cupric complex reduction process as described above.

Gong et al (1993) divided the leaching of gold in a copper-ammonia-thiosulfate system into a two stage reaction.

The first stage is controlled by the interface reaction as described above. The second is a diffusion process of

reaction through a layer of decomposed products. They suggested that the kinetics of leaching gold from an

auriferous pyrite concentrate could be controlled by a corrosion reaction of ammonia-thiosulfate-copper system

on pyrite. However no other workers have reported a corrosion process


19

DISSOLUTION OF SILVER BY THIOSULFATE LEACHING

The Eh-pH diagrams for the ammoniacal thiosulfate and silver system, for high and low concentrations of

ammonia and thiosulfate are shown in Figure 10. Silver combines preferentially with thiosulfate, while silver

ammonia complexes are only stable at low thiosulfate concentrations (Figure 11). Flett et al. (1983) observed

better recovery of silver in the absence of ammonia. Zipperian et al (1988) observed that at higher reagent

concentrations, the concentration of both ammonia and copper (II) in solution affected silver extraction. The Eh-

pH diagram indicates similar potentials are required to leach silver to that required for gold dissolution. If the

potential is too low, silver will remain undissolved or precipitates as a silver sulfide (Ag2S) over the whole pH

range from 0-14. At high potentials and pH values, silver oxide (Ag2O3) precipitates.

The leaching mechanism for silver metal by ammoniacal thiosulfate solutions with copper is similar to that of

gold where an oxidant is necessary (Lukomskaya et al., 1984; Zipperian et al., 1988). However the leaching

mechanism of silver sulfide is reported to occur by the substitution of copper for silver in the sulfide matrix with

silver complexing with thiosulfate in solution as follows (Briones and Lapidus, 1998):

2Cu+ + Ag2S 2Ag

+ + Cu2S (chalcocite) + 4S2O3

2- 2Ag(S2O3)2

3-

Cu2+

+ Ag2S 2Ag+ + CuS (covellite) + 4S2O3

2- 2Ag(S2O3)2

3-

Silver leaching was found to increase with increasing copper and thiosulfate concentrations, and to decrease with

increments of ammonia concentration. This reaction is also controlled by optimising the ammonia/thiosulfate

ratio.

The presence of silver also increases the dissolution of gold at least 6 times more than that of pure gold reacting

with thiosulfate solution (Webster, 1986). In addition, electrum alloys (64 at% Ag) dissolve faster than pure gold.

In terms of leaching kinetics, gold leaches faster than silver. This is most probably associated with the thiosulfate

leaching mechanism being different. In addition, Zipperian et al. (1988) reported silver extraction more sensitive

to changes in reagent concentration than gold extraction.


20

(a)

(b)

Fig 10 Eh-pH diagram for the silver, thiosulfate and ammonia systems (a) High reagent concentrations 5X10-4

m

Ag; 1MS2O32-

;1M NH3/NH4+and (b) low reagent concentrations 5X10

-4m Ag; 0.1MS2O3

2-;0.1M NH3

14121086420

2.0

1.8

1.6

1.4

1.2

1.0

0.8

0.6

0.4

0.2

0.0

-0.2

-0.4

-0.6

-0.8

-1.0

Eh (Volts)

pH

Ag

AgAg2S

Ag2O3

Ag(S2O3)35-

(aq)

Ag(S2O3)23-(aq)

14121086420

2.0

1.8

1.6

1.4

1.2

1.0

0.8

0.6

0.4

0.2

0.0

-0.2

-0.4

-0.6

-0.8

-1.0

Eh (Volts)

pH

Ag

AgAg2S

Ag2O3

Ag(S2O3)35-(aq)

Ag(S2O3)23-(aq)


21

(a)

(b)

Fig 11 Distribution of silver species in the silver, thiosulfate and ammonia systems (5X10-4

m Ag; 0.1MS2O32-

;

0.1M NH3,pH 9.5, Eh 0.250V) with varying (a) ammonia and (b) thiosulfate concentrations.

0.00E+00

1.00E-04

2.00E-04

3.00E-04

4.00E-04

5.00E-04

-6 -5 -4 -3 -2 -1 0 1

Log [NH3] (M)

Co

nc

en

tra

tio

n (

M)

Ag(NH3)2+

Ag(S2O3)23-

Ag(S2O3)35-

0.00E+00

1.00E-04

2.00E-04

3.00E-04

4.00E-04

5.00E-04

-6 -5 -4 -3 -2 -1 0 1

log [S2O32-

] (M)

Co

nc

en

tra

tio

n (

M)

Ag(NH3)2+

Ag(S2O3)35-

Ag(S2O3)23-

Ag(S2O3)-


22

STABILITY OF THIOSULFATE

One of the major problems in thiosulfate leaching is the high consumption of thiosulfate during gold leaching.

High consumption of thiosulfate is mainly caused by its decomposition in solution, or through its loss to tailings

by adsorption onto solids. Zipperian et al., (1988) reported a loss of up to 50% of thiosulfate in ammoniacal

thiosulfate solutions containing copper.

Thiosulfate is a metastable anion that tends to readily undergo chemical decomposition in aqueous solutions. The

factors that influence the stability of thiosulfate are the concentration and pH of solutions, the presence of certain

metals, the presence of sulfur metabolising bacteria, and exposure to ultraviolet light (Dhawale, 1993). Dilute

solutions of thiosulfate (<0.01M) decompose more rapidly than concentrated solutions (>0.1M). Thiosulfate

solutions prepared in freshly boiled, double-distilled water or distilled water are very stable if stored in an air

tight bottle.

Figure 12 presents the Eh-pH diagram for the metastable S-H2O system (Kametani and Aoki, 1976; Osseo-Arare,

1989). The thermodynamically stable species (i.e. HSO4- and SO4

2-) are omitted from consideration to reveal the

metastability domain of species such as thiosulfate (S2O32-

), tetrathionate (S4O62-

) and sulfite (SO3-).

Under alkali conditions, a number of metastable sulfur species, such as sulfite (SO32-

), thiosulfate (S2O32-

),

polythionates (SnO62-

. 2<n<6) and polysulfides (Sn2-

) are found to occur. The sulfur–water system is more

complicated than the metal-water systems because sulfur is mulitvalent and easily forms sulfur chains and

colloidal precipitates. Thermodynamically sulfate is more stable under the leaching conditions.

It can be seen that S2O32-

is located in a narrow elongated stability field in the neutral to basic pH region. To

maintain thiosulfate in solution at ambient conditions, the solution would have to be maintained within this

region. Once formed, the metal thiosulfate complexes are stable over a larger pH-Eh range as observed earlier in

Eh-pH diagrams.

Fig 12 Eh-pH diagram for the metastable S-H2O system {S} = 1.0 M. The S-H2O system without thiosulfate is

superimposed to show an increase in tetrathionate (S4O62-

) domain.

1614121086420-2

2.0

1.8

1.6

1.4

1.2

1.0

0.8

0.6

0.4

0.2

0.0

-0.2

-0.4

-0.6

-0.8

-1.0

S - H2O - System at 25.00 CEh (Volts)

pH

S

S2O62- (aq)

SO32- (aq)

H2S(aq)HS- (aq)

H2SO3 (aq)

S2-

S2O32- (aq)

S2O82- (aq)

S4O62- (aq)


23

Under oxidising conditions, thiosulfate will be oxidised to sulfate or tetrathionate:

S2O32-

+ O2 + 2H2O = 2SO42-

+ 4H+

2S2O32-

+ 1/2O2 + H2O = S4O62-

+ 4OH-

or disproportionates into sulfate and elemental sulfur:

2S2O32-

+ H2O = 2SO42-

+ 4S + OH-

or disproportionates into sulfite and sulfide ion:

3S2O32-

+ 6OH- = 4SO3

2- + 2S

2- + 3H2O

S2O32-

= SO32-

+ S0

depending upon the conditions.

Thermodynamically, thiosulfate oxidises to tetrathionate between a pH range of 4 to 6. Otherwise thiosulfate

oxidises to other sulfur species, such as S2O62-

or SO32-

. In the presence of copper and ammonia, however some

thiosulfate oxidises to tetrathionate at pH 8 to 10 (see Table 3). The oxidative degradation of thiosulfate to

tetrathionate is promoted by the cupric ion (Hemmati et al., 1989, Byerley et al., 1973ab, 1975). Thus the

amount of cupric ion addition, or the concentration of cupric ions is an important factor in thiosulfate stability

and reagent management.

Under non-oxidising conditions some thiosulfate (~62.5%) can be regenerated slowly from the decomposition of

tetrathionate to higher or lower polythionate ions through the formation of trithionate ions (Naito et al., 1970;

Byerley et al., 1973). The reaction is highly catalysed by the presence of thiosulfate ions according to the

reaction scheme:

S4O62-

+ S2O32-

S5O62-

+ SO32-

SO32-

+ S4O62-

S3O62-

+ S2O32-

S5O62-

+ 3OH- 5/2S2O3

2- + 3/2H2O

---------------------------------------------------------------

2S4O62-

+ 3OH- 5/2S2O3

2- + S3O6

2- + 3/2H2O

In ammoniacal thiosulfate leaching of gold, increasing the concentration of thiosulfate in solution results in an

increase in thiosulfate consumption (Cao et al., 1992). Efforts have been made to reduce consumption by

reducing the concentration of thiosulfate in solution and leaching for extended periods (e.g. 24 hours) (Cao et al.,

1992). Alternatively, sulfite and sulfate have been added to the leaching solution to stabilise the thiosulfate as

described below.

Effect of sulfite

The presence of sulfite inhibits the decomposition of thiosulfate (Kerley , 1981,1983).

An equilibrium reaction occurs in the thiosulfate liquor as represented by the following equation (Hemmati et al.,

1989)

SO32-

+ S0 = 3S2O3

2-

The presence of sulfite ions is claimed to prevent the formation of any free sulfide ion and the precipitation of

gold or silver from solution. Maintaining a level of 0.05% has been found to stabilise thiosulfate (Kerley, 1983)

but this will also lower the Eh of the solution and reduce Cu(II) in solution. It has also been shown that sulfite is

oxidised by copper (II) ion to sulfate and dithionate depending upon the reaction conditions

Manganese containing ores have a high requirement for sulfite ion because of the oxidising capability of various

manganese compounds (Kerley, 1981). In addition high concentrations of oxygen and/or copper in solution will

decompose sulfite to sulfate. Kerley (1983) observed that the addition of sulfite to reduce strongly oxidising

mineral components in the ore increased gold recovery from about 5.8% to 84.5%.


24

Effect of sulfate

Jiexue and Qian (1991) and Hu and Gong (1991) claimed that gold could be extracted from sulfide ores with

thiosulfate-sulfate instead of thiosulfate-sulfite solutions. They found that the loss of thiosulfate could be reduced

and the gold extraction increased by adding sulfate to the solution. They suggested that the following reaction

takes place:

SO42-

+ S2-

+ H2O = S2O32-

+ 2OH-

However this is most unlikely as sulfate is very stable.

Sulfate ions are unreactive in ammoniacal thiosulfate leaching system, but are also a decomposition product of

thiosulfate. A build up adversely affects the activity of the solution. The addition of lime has been recommended

if such a buildup of sulfate does occur. The lime reacts with the sulfate ions to precipitate gypsum.

EFFECT OF OTHER CATIONS IN SOLUTIONS

In general, leaching gold by thiosulfate permits a decreased interference from foreign cations in comparison with

the conventional cyanidation method. However, Perez (1987) reported that at pH levels lower than 8.0, metallic

iron (from grinding media) and iron salts dissolve in solution resulting in a decrease in gold dissolution. The

ferric ion is reduced to a ferrous ion and the thiosulfate ions are oxidised to tetrathionate ions (Perez and Galaviz,

1987). Tetrathionate has no lixivating action on gold or silver.

The dissolution of copper present in the ore can be used directly in the leaching process as described above.

Moderate concentrations of cobalt, nickel, and manganese can also be in solution at temperatures above

ambience in ammonium thiosulfate solutions. However, at pH values greater than 10 they are generally in low

concentrations (Niinae et al, 1996).

PASSIVATION OF GOLD

Early work at higher concentrations of reagents reported the precipitation of copper sulfide and the promotion of

the oxidation of thiosulfate to tetrathionate. Clearly, the formation of CuS and sulfur layers on gold particles will

hinder their dissolution.

Bagdasaryan et al. (1983) and Pedraza et al. (1988) observed a sulfur layer as well as copper sulfide in a

thiosulfate-copper sulfate system. Both elemental and sulfide sulfur can be provided by the decomposition of

thiosulfate in alkaline solution.

Electrochemical impedence spectral studies have shown that gold passivation can occur in the absence of copper

(MacDonald, 1990). The appearance of an inductive arc in the potential range of gold dissolution in sodium

thiosulfate solution is assigned to the passivation of the gold electrode accompanying its dissolution. The belief

is that elemental sulfur may be formed preventing thiosulfate from diffusing to the gold surface and hence,

inhibiting the gold dissolution. The elemental sulfur coating is formed either by the adsorption of elemental

sulfur, or by the oxidation of sulfide ion on the gold surface.

The passivation of gold can be prevented with the presence of ammonia, adjustment of Eh-pH conditions and the

presence of O2 as described earlier.

EFFECT OF TEMPERATURE ON THIOSULFATE LEACHING

Higher gold and silver dissolution rates have been reported at temperatures in the range of 40-60C (Berezowsky

and Weir, 1989; Qian and Jiexue, 1989; Hemmati et al., 1989;Tozawa et al., 1981; Zipperian et al., 1988). In

contrast, Abbruzzese et al. (1995) found in their studies that increasing temperature decreased gold recoveries by

20%. They ascribed this result to passivation by cupric sulfide formed by the thermal reaction between Cu(II)

ions and thiosulfate as follows :

Cu2+

+ S2O32-

+ H2O = CuS + SO42-

+ OH-


25

However, the decrease may have been caused by loss of ammonia from solution at higher temperature

destabilising the Eh-pH leaching conditions. Such problems may have been overcome by the addition of sulfite

as described above.

Temperatures above 60C make it difficult to maintain ammonium hydroxide in solution (Kerley, 1981). In

addition heating a large quantity of material would not be practical unless the ore had been previously treated in

some way at elevated temperatures.

AUTOCLAVE LEACHING OF GOLD

To overcome problems associated with passivation of gold through the formation of sulfur and copper sulfide

layers on gold particles during thiosulfate leaching, most work in the early days was carried out at elevated

temperatures and pressures in autoclaves. At temperatures above 100C the layer of elemental sulfur that is

formed by the oxidation of thiosulfate can be redissolved:

4S + 6OH- S2O3

2- + 2S

2- + 3H2O

Thiosulfate is regenerated in alkaline solutions, whereas sulfur is reprecipitated in acidic solutions. Ter-

Arakelyan et al. (1984) noted that ammonium sulfate promotes the oxidation and dissolution of the sulfide layer

on the surface of gold only at elevated temperatures and oxygen pressures.

LEACHING WITH POLYSULFIDE

In addition to thiosulfate, polysulfides will also be present during leaching. These can be produced from the

disproportionation of elemental sulfur or thiosulfate decomposition.

Investigations into the use of polysulfides to leach gold has been carried out by Chen et al., (1996). From

electrochemical studies carried out on a gold electrode in polysulfide solutions, gold dissolution occurred as a

result of absorption of polysulfide on its surface, accompanied by the oxidation of polysulfide as follows:

Anode: Au/Sx2-

AuS- + (x-1)S

0 + e

-

Au/Sx2-

Au/Sx + 2e-

Cathode: S0 +2e

- S

2-

The cathode reaction indicates that the elemental sulfur acts as an oxidant for gold dissolution. Thus gold can be

leached in polysulfide solution without the addition of any oxidant. Chen et al. (1996) reported at 50C, 90%

gold extraction from a sulfide concentrate without the addition of an oxidant. However, a relatively high

polysulfide concentration (>2M) is required for high gold extraction.

In mixtures of thiosulfate and polysulfides, the polysulfides only act as a lixivant when no oxidant is present. In

the presence of copper, which is required as the oxidant in thiosulfate leaching, polysulfides precipitate with

copper to form CuS.

GENERATION OF THIOSULFATE IN SITU

As thiosulfate systems can be economically unacceptable due to high reagent consumption, several authors have

investigated the possibility of generating thiosulfate in-situ.

Thiosulfate and polysulfides can be produced by the disproportion of elemental sulfur with OH- ions in

ammoniacal solution through a phase transfer catalysis (Deng et al., 1984) or from hydrated lime solution with

sulfur (Zhang et al., 1993) as follows

4 S0 + OH

- S2O3

2- + 2S

2- + 3H2O

Sulfide ions are then able to form polysulfides with elemental sulfur as follows

S2-

+ S0x-1 = Sx

2-.


26

and dissolve gold as described above.

However, thiosulfate degrades with time to sulfate and sulfide.

Zhang et al (1993) using the Lime Sulfur Synthetic Solution (LSSS) method obtained gold recoveries over a two

stage process of 99% for a high lead sulfide concentrate and nickel-copper matte residue respectively.

Thiosulfate can also be produced by the reaction of hydrogen sulfide on aqueous solutions of sulfites:

2HS- + 4HSO3

- = 3S2O3

2- + 3H2O

or by boiling aqueous solutions of metal sulfite with elemental sulfur

Na2SO3 + S = Na2S2O3

In this method, a slurry of sodium sulfite is digested with an excess of sulfur to produce thiosulfate. The sodium

sulfite slurry is prepared by passing sulfur dioxide through soda ash solution.

Oxidation of polysulfides generates thiosulfate as follows (Greenwood et al, 1984):

Na2S5 + 3/2 O2 = Na2S2O3 + 3S

Kerley (1983) described a method of generating thiosulfate by the following reaction:

2NH3 + SO2 + S + H2O (NH4)2S2O3

Ammonia can be recycled while elemental sulfur and sulfur dioxide gas is added in sufficient amounts to make

thiosulfate. The reaction proceeds very slowly and so an excess of sulfur is required. Free ammonia should

always be present to avoid acid conditions. Sulfur dioxide gas is readily dissolved in water at pH<2. By

increasing the pH, sulfur dioxide forms sulfite as follows

2NH3

SO2(aq) + H2O = H+ + HSO3

2- H2O + SO3

2- + 2NH4

+

In the presence of sulfur, thiosulfate is produced as follows

3H2O + 4SO32-

+2S = 3S2O32-

+6OH-

Using this method Kerley (1983) was able to recovery 90% silver from a 16oz/t Ag grade ore containing 10.5%

manganese.

Thiosulfate formation and complexing with gold has been observed in natural and particularly in supergene

enivronments (Bowell et al., 1993; Benedetti and Boulegue, 1991; Kucha et al., 1994). In alkaline solutions,

pyrite oxidation proceeds through the intermediate tetrathionate ion (at pH 6-7) and thiosulfate ions (pH >7).

Goldhaber (1983) showed that sulfate is directly formed by pyrite oxidation only at pH < 6. The breakdown of

meta-stable thiosulfate to sulfate is delayed by an ‘induction period’, the duration of which increases with

alkalinity (Rolla and Chakrabarti, 1982). However, the concentration of thiosulfate in solutions is low. No doubt

certain sulfur-oxidising enzymes and bacteria may increase thiosulfate concentrations initially (Lakin et al.,

1974, Webster, 1986).

Cao et al (1992) tried to generate thiosulfate in-situ by leaching a copper-gold ore in ammonia solution. No

thiosulfate was present but up to 22% gold was recovered as a result of gold-copper-ammonia complexes being

formed.

It has been reported that significant amounts of thiosulfate can form during ammoniacal pressure oxidation

leaching of complex copper-nickel-cobalt sulfide concentrates (Forward et al., 1955).


27

DECOMPOSITION OF SULFIDES IN THIOSULFATE LEACH SOLUTION

From X-ray powder diffraction studies Qian and Cao (1993) observed less pyrite present in thiosulfate leached

residues than in the original ore. They concluded that a ammonia-thiosulfate solution containing copper, leaches

pyrite by a corrosive process. Scanning electron microscopy showed clear scarring on pyrite leached grains.

These authors also observed some dissolution of chalcopyrite in the residues.

Copper sulfide minerals, other than chalcopyrite, also dissolve readily in aerated thiosulfate leach solution,

particularly when ammonia is in solution.

THIOSULFATE LEACHING OF ORES

Table 4 shows a summary from selected literature of the various thiosulfate leaching conditions that have been

used. After initial studies around 1980, renewed interest occurred in the 1990’s. Most work has been carried out

on complex ores containing high copper, carbonaceous ores or ores containing high concentrations of lead, zinc

or manganese. The gold dissolution rates and percentage recovery will vary depending on the deportment of gold

in the ores. A wide range of conditions appear to have been used, including relatively severe conditions,

compared with those used in the cyanidation process.

The gold concentration in the ore has varied from between 1 and 62 g/ton. The addition of copper has varied

from 0.001 to 0.1 mol/L, whereas ammonia concentration has varied from 0.1 to 6 mol/L and thiosulfate

concentration from 0.1 to 2 mol/L. The pH conditions are all alkaline with most leaches carried out at near pH 9-

10. Generally the leaching time has been over a relatively short period of a few hours. Reagent consumption

appears to be lower where reagent concentrations are low and the ore leached over longer periods of time.

The patents taken out that refer to thiosulfate leaching are listed in Table 5. With the exception of Yen et al.

(1998), recent papers and patents use much more dilute reagent concentrations and low Cu2+

concentrations in

order to minimise thiosulfate oxidation.


28

TABLE 4 List of conditions used in literature

Ore type Gold (g/ton) Temperature pO2 (kPa) Retention (hrs) S2O32-

(M) NH3 (M) Cu2+

(M) SO32-

(M) pH %Recovery Consumption

Tozawa et al , 1981 Gold plates 99.99% 65 100 3 0.5 1 0.04 - -

Kerley, 1981,1983 sulfide 18% 2% 4g/L 2% 7-9 95 4Kg/t

2% Mn

Block-Bolten and Torma Zn-Pb sulfide flotation 1.75 21-50 air 2L/min 1 0.125-0.5 0.75 - - 7-9 90 45 Ib/t

1985, 1986 22.5g/t Ag,0.7%Zn,0.5% Pb

Zipperian et al , 1988 Rhyolite ore 3 50 atm 2 2 4 0.1 - 10 90% 50%

7g/Kg MnO2

Hemmati et al , 1989 Carbonaceous 14.74 35 103 4 0.71 3 0.15 0.22 10.5 73% 15-19%

2.5% org C

Caixia and Qiang, 1991 Oxidised ore 4.78 30-65 atm 2 1-22% 1.36-8.86% 0.05-2% 1% - 93.9 40 Kg/t

0.05% Cu,

Hu and Gong, 1991 0.048% MnO2 50.4 40 1L/min 1-2 1 2 0.016 0 - 95.6 -

3.19% Cu

Murthy, 1991 Pb-Zn sulfide 1.75 21-70 atm 3 0.125-0.5 1 - - 6.9-8.5 95% -

22.5g/tAg, 0.44% Cu, 0.68% Zn, 0.54% Pb

Cao et al. , 1992 Sulfide conc 62 60 atm 1-2 0.2-0.3 2-4 0.047 - 10-10.5 95 4.8 Kg/t

3% Cu

Langhans et al, 1992 Oxidised ore 1.65 ambient atm 48 0.2 0.09 0.001 - 11 90 2kg/t

0.02% Cu

Wan et al , 1994 Carbonaceous 2.4 ambient atm 12-25 days 0.1-0.2 0.1 60ppm - 9.2-10

1.4% C, 1.0% S

Abbruzzese et al , 1995 Gold ore 51.6 25 atm 3 2 4 0.1 - 8.5-10.5 80

Groudev et al , 1996 Bacteria leached ore 3.2 ambient atm 15g/l added to pH 9.5 0.5g/l 0.5g/l 9.5-10.0 80

15.2 g/t,0.14%Cu,0.91%S

Marchbank et al , 1996 Sulfide carbonaceous 3-7 55 atm 4 0.02-0.1 2000ppm 500ppm 0.01-.05 7-8.7 70-85

Yen et al ,1996 Gold-copper ore 7.26 ambient atm 24 0.4 0.2 0.03 - 11 90

1.4g/t Ag, 0.3% Cu

Wan, 1997 Sulfide carbonaceous 1-3 ambient atm -

Wan and Brierley, 1997 Carbonaceous sulfide 1-3 ambient atm 91-116days 0.1 0.1 0.005 - 9 50.7-65.7 5.2-8.4

Yen et al , 1998,1999 Gold copper ores 7.2-7.9 ambient atm 24 0.5 6 0.1 - 10 95-97 30kg/t

~0.36% Cu

Thomas et al , 1998 Pressure oxidised sulfide ore 2.5 45-55 atm 12 0.03-0.05 ~500-1000ppm 10-100ppm 0.01-0.05 7.5-7.7 80-85


29

TABLE 5 Patents on thiosulfate leaching

Authors Title Patent No

Kerley, Jr 1981 Recovery of precious metals from difficult ores US 4,269,622

Kerley, Jr, 1983 Recovery of precious metals from difficult ores US 4,369,061

Perez and Galaviz, 1987 Method for recovery of precious metals from difficult ores

with copper-ammonium thiosulfate

US 4,654,078

Genik-Sas-Berezowsky;

et al 1978

Recovery of precious metals from metal sulfides US 4,070,182:

Wan et al., 1994 Hydrometallurgical process for the recovery of precious

metals values from precious metal ores with thiosulfate

lixiviant

US 5,354,359

Marchbank et al., 1996 Gold recovery from refractory carbonaceous ores by

pressure oxidation and thiosulfate leaching

US 5,536,297

Pappas (1997) Process for recovering gold from oxide-based refractory

ores

UK GB 2,310,424 A

Thomas et al., 1998 Gold recovery from refractory carbonaceous ores by

pressure oxidation, thiosulfate leaching and resin-in-pulp

US 5,785,736

Leaching copper containing ore

The presence of reactive copper in auriferous ore is well known to cause high reagent consumption

using the conventional cyanide method. The advantages of using ammoniacal thiosulfate containing

copper to treat such ore has been investigated. Genik-Sas-Berekowski (1978) patented a technique for

the use of ammonia thiosulfate as a secondary leach for the recovery of silver and gold in conjunction

with a hydrometallurgical process for the recovery of copper from copper bearing sulfide ores.

Chalcopyrite ores containing gold were leached in ammonium sulfate with an O2 over-pressure of 20

psi before treatment with 65-130 g/L of ammonium thiosulfate within pH values of 8.5 to 9.5. In some

cases high gold and silver recoveries (90%) were observed. However, the method did not take into

account the instability of thiosulfate and the time related instability causing loss of recovery of gold.

Kerley (1981, 1983) described a process in which ores which are difficult-to-treat by the conventional

method were treated by lixiviation in ammonia thiosulfate solutions containing copper to extract gold

and silver. Sulfite ions were provided by adding ammonium sulfite or ammonium bisulfite to the

leaching solution in order to minimise the oxidation of thiosulfate. The pH recommended was at least

7.5 or higher.

Perez and Galaviz (1987) modified Kerley’s process to inhibit Fe3+

oxidation of thiosulfate (and

possibly Cu2+

cementation) caused by the dissolution of metallic iron from grinding media present in

the treated ore. Their process involved treating copper containing gold ore with copper ammonium

thiosulfate in which the pH of the lixiviating solution was maintained at a minimum of 9.5. Copper

cement was then used in a subsequent precipitation process on which the gold and silver was

precipitated without causing precipitation of copper from the lixiviant solution. Trial runs on a pilot

plant scale were carried out at LaColorada, in the State of Senora, Mexico.

The pilot plant layout that existed at LaColorada is shown in Figure 13. The ore was ground to increase

the amount of surface area of the ore exposed to the lixiviating solution. Before milling, copper

ammonium thiosulfate and water were added to the ore until the concentration required was obtained.

Anhydrous ammonia was added to the copper ammonium thiosulfate to maintain the pH at a level of at

least 9.5. After grinding to approximately 200 mesh, the viscosity of the slurry was lowered by adding

more water and ammonium thiosulfate to bring the solids in the slurry to 40%. This also lowered the

pH to around 8 or 9. An agitator step was then performed to increase reagent contact with gold and

silver. After agitating for about 1.5 hours, the slurry went to a thickening tank and the pregnant solution

was removed to another agitator tank. Copper cement from a container was fed into the agitator tank to

precipitate the gold and silver from solution which settled in the settling tank. The gold and silver

precipitate was then collected and the spent solution reused. High gold recoveries were obtained, but

there is no mention of reagent consumption.


30

Fig 13 Flow chart of pilot plant layout that existed at LaColorada, in the State of Senora, Mexico

(after Perez and Galaviz, 1987).

Cao et al (1992) reported a low thiosulfate but high ammonia concentration leaching system that has

the potential to be economic for treating sulfide gold ores. Over 95% gold was extracted from

auriferous sulfide ore containing high copper (62g/t Au and 3.1% Cu). The conditions used were 0.2-

Precious

Metal

Precipitate

Settling

Tank

Agitator

Tank Copper

cement

SO2

to

lower

pH

Tailings

Filtration

Thickener

Agitator

Tank

Ball Mill

Fine ore

Bin

NH3

Cu(NH3)42+

(NH4)2(S2O3)

Pregnant

solution

Water + (NH4)2S2O3


31

0.3M (NH4)2S2O3, 3 g/L Cu2+

, 2-4M NH4OH, 0.5-0.8M (NH4)2SO4, at 60C for 1-2 hours aeration.

However this work has only been carried out under laboratory conditions, and losses of thiosulfate and

ammonia were not reported.

An evaluation of the potential of heap leaching, or in-situ copper catalysed thiosulfate leaching of low

grade ore was carried out by Langhans et al (1992). The work concentrated on investigating the

feasibility of using thiosulfate under long term leaching conditions and at lower reagent concentrations.

From 48 hour leaching tests they obtained 83% gold extraction with a consumption of 0.4 kg/t of

thiosulfate at 0.2M S2O32-

, 0.00625M SO32-

, 0.001M Cu(II) and 0.09M NH3. However the leaching

studies were conducted on low grade oxidised ore with traces of sulfides.

Preg robbing type ores

The presence of carbon in an ‘active’ form will adsorb gold that has dissolved in the conventional

cyanide process, thus robbing the lixiviant. This process is known as ‘preg-robbing’. By virtue of the

fact that gold thiosulfate adsorption on to carbon is low, thiosulfate leaching of carbonaceous ores is

potentially economical, therefore attracting significant industry research.

Hemmati et al. (1989) investigated the thiosulfate leaching of gold from carbonaceous ore that

contained 2.5% organic carbon. The optimum conditions for gold extraction were found to be 35C and

103 kPa oxygen over pressure, pH 10.5, 3M NH3 with a lixiviant containing 0.71M (NH4)2S2O3, 0.15M

CuSO4 and 0.1M (NH4)2SO3. Extraction of gold was 73% in 4 hours compared with only 10%

observed using cyanide over 24 hours.

In the case of low-grade carbonaceous ores, they can be processed without pretreatment by heap

leaching with thiosulfate. However, ores containing high sulfide concentrations require at least partial

oxidation prior to thiosulfate leaching.

Newmont Gold Co patented a microbial innoculation/agglomeration process for rapid initiation of bio-

oxidation of refractory ores in heaps followed by thiosulfate leaching to avoid the preg-robbing activity

of carbonaceous matter (Brierley and Hill, 1993, Wan et al., 1994; Wan and Brierley, 1997). The work

was carried out in trials on both laboratory and pilot plant scale experiments. The process consists of

growing a dense cell suspension of the naturally occuring iron-oxidising bacteria Thiobacillus

ferrooxidans and Leptospirillum ferrooxidans and then adding this bacterial culture to the ore as the

heap is formed. The distribution of bacteria throughout the heap is achieved with noticeable

agglomeration of the fine ore particles. The effluents from the heaps are collected in either a tank or

pond and recirculated to the top of the heap. The progress of the biooxidation pretreatment process is

monitored by measuring pH, Eh and Fe2+

/Fe3+

ions in solution. After bio-oxidation, an agglomeration

/neutralisation step is carried out to neutralise the acidic ore for leaching with thiosulfate under alkaline

conditions. This is achieved by washing the heap in water to displace soluble acid and iron salts. The

washed heap is further neutralised by blending in lime or soda ash.

A schematic diagram of the heap leach process used by Newmont Gold Co during the thiosulfate

leaching stage is shown Figure 14. After passing through the heap, the pregnant lixiviant solution is

recovered at the bottom of the heap and recirculated, either continuously or intermittently. After the

precious metals values are recovered from the lixiviant solution by precipitation, the solution is

recirculated to the static heap. Using this method Wan et al., (1994) obtained gold recoveries as high as

70% whereas the conventional cyanidation process resulted in only about 20% (Wan and Brierley,

1997). They recommended a leach solution with pH between 9.2-10.0, 0.1-0.2M (NH4)2S2O3 and/or

Na2S2O3, at least 0.1M ammonia and up to 60ppm of Cu(NH3)42+

to act as a catalyst for gold oxidation.


32

Fig 14 A schematic diagram of a gold and silver leaching and recovery process for static heaps using

thiosulfate (after Wan et al., 1994).

A similar process was also developed by Pappas (1997) for treating carbonaceous oxide ores.

An alternative to bacterial oxidation as a pretreatment of sulfide rich carbonaceous ores, is to fine mill

the ore prior to thiosulfate leaching (Wan et al., 1994). A schematic diagram illustrating such a process

is shown in Figure 15. Finely ground carbonaceous ore is initially slurried with water, thiosulfate,

copper sulfate and ammonia in a slurry preparation unit. Each reagent is added in appropriate quantities

to obtain the right pH and oxidising concentration conducive to gold extraction. During the process,

stripped thiosulfate lixiviant is pumped from the precious metal recovery unit into the slurry

preparation unit thus reducing the amount of extra reagents and water required to maintain the desired

conditions for leaching. The ground slurry is transferred to a heat exchanger where the temperature can

be increased to about 45ºC to improve gold extraction. The slurry is then passed to a stirred tank

reactor where the extraction of gold by the thiosulfate lixiviant takes place. After the appropriate

contact interval, lixiviated ore slurry is transferred from the stirred tank reactor to a conventional

separator where the separator overflow is the pregnant thiosulfate lixiviant and the underflow is the

leached residue. The leached residue is transferred to the tailings and the precious metal in the pregnant

solution is recovered by means of precipitation. The retention time for the whole process is governed

by the composition of the ore and its particle size.

Heap

Pregnant

Lixiviant

Solution

Storage

Ammonium

Thiosulfate

Precious

Metal

Recovery

system

Ammonium

Thiosulfate

NH3

Cu2+

Pump


33

Fig 15 Major steps in a gold and silver leaching and recovery process for finely ground ore material

using thiosulfate (after Wan et al., 1994).

Barrick Co. have also patented a process for recovering gold from an ore having a combination of

refractory sulfide and carbonaceous material using pressure oxidation followed by thiosulfate leaching

(Marchbank et al., 1996). The oxidised ore slurry was mixed with thiosulfate solution maintained at a

pH between 7 and 8.7. The preferred lixiviant was ammonium thiosulfate (0.025-0.1M) with a

residence time of 60-240 minutes at temperatures between 45-55C. The cupric ion concentration was

kept in the range of 5-50ppm. Sufficient ammonia was added to maintain the copper in solution with a

minimum molar ratio of 4:1 NH3:Cu. To reduce thiosulfate consumption, 0.01-0.05M sulfite was

added in the form of sodium metabisulfite or through additions of sulfur dioxide gas. About 70-75%

gold was extracted from the ore.

Recently the Barrick Gold Corporation patented a combined pressure oxidation, thiosulfate and resin-

in-pulp process for treatment of refractory gold ores (Figure 16) (Thomas et al., 1998). In this process

ore is ground to a size with about 65-85% passing 200 mesh and thickened to about 40-50% solids.

Sodium carbonate is added to ensure that pressure oxidation is carried out under alkaline conditions

and Cl- is added to improve the kinetics and to facilitate oxidation. The ore which is pressure oxidised,

leaves the autoclave at about 35% solids and is directed to a leaching operation where it is contacted

with ammonium thiosulfate (5g/L) and copper sulfate (25ppm Cu). The slurry of gold-bearing leachate

and solid residue leaving the leaching circuit contains in the range of 1-5ppm gold and is directed to a

Resin-in-pulp (RIP) circuit where gold and copper are loaded onto a strong base resin to about 1-5Kg/t

Au and about 10-25 kg/t Cu. Copper is eluted from the resin using ammonia thiosulfate (200g/L) and

gold is eluted using potassium thiocyanate (200g/L). The copper-bearing eluate is returned to the

leaching circuit while the gold eluate is either electrowon or precipitated. Clearly the resin is not

selective for gold, and loads both Au and Cu in proportion to their concentration in solution. This

approach is therefore applicable only to copper-free ores using low levels of Cu(NH3)42+

as a catalyst.

Slurry

Preparation

NH3 Water

Copper

sulfate

Thiosulfate

Heat

Exchanger

Stirred

Tank

Reactor Separator

Precious

Metal

Recovery

Ground

Carbonaceous ore

Sulfur

Pretreatment


34

Fig 16 Flow sheet illustrating the pressure oxidation, thiosulfate and resin in pulp recovery process for

refractory gold ores (after Thomas et al., 1998)

Manganese containing ore

Kerley (1981,1983) used ammoniacal thiosulfate to leach gold and silver from an ore containing 2%

manganese. He used a solution containing 18% (NH4)2S2O3, 3% (NH4)2SO3, 2% NH4OH and 4g/L Cu.

Although gold and silver recoveries were high, reagent consumption was 4 kg/t (NH4)2 S2O3 , 3 kg/t

(NH4)2SO3 and 0.5 kg/t Cu. Manganese type ores containing precious metal ores have a high

consumption rate of sulfite with the following reaction taking place

MnO2 + 2(NH4)2SO3 + 2H2O MnS2O6 + 4NH4OH

Ammonium

Thiosulfate

Tailings

Disposal

Liquid/

Solid

Separation

optional

Liquid

Water Solids

Water

Treatment

Grinding

Ore

Autoclave

Pressure

Oxidation

Leach

RIP

Cu Elution

Au Elution

Electrowinning

or Precipitation

Potassium

Thiocyanate

Cu/(NH4)S2O3

Resin

CuSO4

(NH4)2S2O3 NH4+

Steam

O2 H2O

NAOH/NACL

or

Na2CO3/NaCl


35

Lead- Zinc sulfide ore

Gold and silver extraction from low grade lead and zinc sulfide flotation tailings was investigated by

Murthy (1991). Thiosulfate leaching at 50 resulted in 95% gold extraction. Block-Bolten and Torma

(1985, 1986) obtained high gold recovery (90%) from a lead-zinc sulfide flotation tailings using a two

step counter current leaching process. The counter current leaching process consisted of the fresh

tailings being in contact with pre-used leach solution and the pre-leached tailings receiving the fresh

leachant. Compared with cyanide heap leaching, the thiosulfate process required only about 1 hour to

achieve 90% gold recovery using up to 0.5M thiosulfate whereas cyanide required at least 24 hours to

extract 50% of the gold. While results are favourable for thiosulfate leaching, experiments were carried

out on a laboratory scale and high gold extraction required a temperature of 50 °C.


36

RECOVERY OF GOLD FROM THIOSULFATE LEACHED ORE

Most research has been concerned with the extraction of gold with thiosulfate and only limited work

has been carried out on the process of recovering gold from solution. In general precipitation methods

have been considered for clarified leach solutions particularly heap leaching, whilst carbon and resins

have been considered for adsorption from slurries.

Precipitation methods

Gold and silver in copper-ammonia thiosulfate solutions can be precipitated out of clarified solution by

the addition of finely divided copper metal (Perez and Galaviz, 1987), zinc metal (Berezowsky and

Sefton, 1979; Panayoto et al., 1994), iron metal (Jiexue and Qian 1989), aluminium metal or soluble

sulfides (Kerley, 1983).

The reduction of metals from solution is a result of charge-transfer reactions. The reaction of gold

recovery by zinc cementation in the presence of ammonia can be represented in the following

reactions:

2 Au(S2O3)23-

+ 2Zn0 + 4NH3 = 2 Au

o + 2 S2O3

2- + Zn(S2O3)2

2- + Zn(NH3)4

2+

Berezowsky and Sefton (1979) observed high consumption of zinc by cupric ion in the thiosulfate

solution and appreciable amounts of zinc and copper in the cement product.

The reaction for the precipitation by zinc can be expressed in the following reaction:

Cu(S2O3)22-

+ Zn0 = Zn(S2O3)2

2- +Cu

0

Increasing the copper ions (such as copper sulfate) in solution can reduce the amount of unreacted zinc

remaining with the gold (Wan et al., 1993). However, cementation of copper results in low grades of

gold. A suitable reductant such as sulfur dioxide can be used to reduce the cupric form to cuprous prior

to zinc dust cementation. This reduces the amount of copper co-precipitating with gold.

In the Merrill-Crowe zinc cementation process used in cyanide lixiviation systems, de-aeration of the

filtered pregnant solution prior to cementation is one of the most important factors for efficient gold

recovery. The presence of oxygen passivates the surface of the zinc and also causes re-dissolution of

the gold precipitate. This results in an excessive consumption of zinc and incomplete recovery of gold.

In this application the presence of copper as well as air results in high zinc consumption. Furthermore

significant dissolved gold is left in the filtered tailings.

Jiexue and Qian (1989) claimed that the use of iron powder prevented the problems observed with the

precipitation of gold using the zinc method. However, it is expected that copper would also precipitate

with iron.

Detailed kinetic studies by Guerra and Dreisinger (1999) observed that increased temperature (30-

50C) and a higher pH/ammonia concentration enhanced cementation performance, whereas the

presence of sulfite and copper ions in solution negatively affected cementation performance. In

addition they observed that measurement of changes in solution potential cannot be used to indicate the

progress of the gold cementation reaction since measured solution potential is dominated by the mixed

potential of side-reactions.

Sodium borohydride can also be used as an efficient agent for reducing gold and silver in clarified

acidic solutions of thiosulfate at room temperature. The Au(I) ion is reduced to metallic gold in the

form of very fine crystals. A complete reduction of gold in thiosulfate solutions can occur with a

sodium borohydride to gold molar ratio of 0.625 at a pH of 6 over a one hour time period (Awadalla

and Ritcey, 1991,1993; Groves and Blackman 1995). However, the presence of ferrous ion, cobalt,

nickel or in particular, copper in solution decrease the efficiency of borohydride to reduce gold because

of extensive co-precipitation of other metals. Most work has been adapted to thiourea leaching rather

than thiosulfate.


37

Gold recovery has also been achieved by sparging or pressurising gold thiosulfate solutions with

hydrogen (Deschenes and Ritcey, 1990). It is inferred, although no evidence was given, that thiosulfate

is not lost.

Finally dissolved gold and silver has been recovered by the addition of a sulfide solution with

regeneration of thiosulfate (Kerley, 1981; Flett et al., 1983):

Ag2S2O3 + (NH4)2S Ag2S + (NH4)2S2O3

Solvent extraction techniques

In general solvent extraction can only be applied to clarified solutions containing relatively high

concentrations of gold and silver. Nevertheless there have been significant efforts at finding suitable

extractants.

Alkyl phosphorus esters

Solvent extraction with alkyl phosphorus esters on gold thiosulfate solutions show high gold recoveries

in alkaline conditions with increased gold recoveries being observed with increased concentrations of

thiosulfate (Zhao et al., 1997). In sodium thiosulfate solution, the extraction reaction using tributyl

phosphate (TBP) can be represented as:

iNa+ + 2Au(S2O3)2

3- + OH

- + mTBP = NaiAu2(S2O3)2-3(OH).mTBP

Where m=1.5-2.5

Recovery of gold was lower with aromatic hydrocarbons than with aliphatic hydrocarbon diluents.

The presence of ammonia and higher concentrations of TBP improved gold recoveries due to the

formation of hydrogen bonds between ammonia and TBP (iNaiAu2(S2O3)3(OH)NH3)2.nTBP, where

n=6-9).

Primary, secondary and tertiary amines

Primary amine (R1R2CHNH2. where R1+R2 = C18-22), secondary amine DNA (dionylamine) and tertiary

amine TOA (trioctylamine) were tested for the extraction of gold from thiosulfate solutions (Zhao et

al., 1998a,b). The extraction capacity decreased in the order of primary>secondary>tertiary amine. It is

suggested that the species extracted by amines from alkaline solutions is different to that from acid

solutions. The extraction reaction of Au(S2O3)22-

by primary amines is proposed as follows:

Au(S2O3)23-

+ 3H+ + 3RNH2 = (RNH3)3Au(S2O3)2

and in ammoniacal thiosulfate solution:

NH4+ + Au(S2O3)2

3- + 2RNH3

+ = NH4(RNH3)2Au(S2O3)2

In general the separation of gold from silver, copper, zinc and nickel in thiosulfate solutions by primary

amines is difficult. However, the addition of ammonia into the thiosulfate solution increased the

difference between the extraction of gold and other metals from the solution leading to good

separation.

Activated carbon adsorption

Activated carbon is preferred for gold recovery from cyanide solutions because it can be added to the

pulp and avoids soluble losses in tailings. Unfortunately it has a low affinity for Au(S2O3)23-

(Gallagher

et al., 1989).


38

Jiexue and Qian (1989) recorded a 30% recovery of gold thiosulfate on activated carbon. For

thiosulfate leached slurries containing up to 16 mg/L of gold, Abbruzzese et al., (1995) observed an

increase in gold recovery from 22% to 43% when the carbon concentration was increased from 5g/L to

60g/L. They also claimed to have achieved a 95% gold recovery after six hours. However increasing

the carbon concentration (120 g/L) resulted in a decrease in gold recovery. Yen et al. (1998) observed

similar gold recoveries of 95% in thiosulfate solutions containing 60g/L of carbon at pH 11 after 6

hours. However the gold loading on carbon at these concentrations was too low to be considered

practical or economic.

The reason for the low loading of gold onto carbon from thiosulfate system is uncertain. It has been

suggested that it could be related to the relatively high negative charge of the complex, steric

limitations due to its molecular structure, or specific interactions of the ligand group with carbon active

sites.

The potentials for the reduction of five gold complexes to gold metal from a solution containing gold at

10-5

M given by McDougall and Fleming (1987) are shown in Table 6.

TABLE 6 Potentials for the reduction of gold complexes to gold metal from a solution

containing gold at 10-5

M

Ligand Ligand E0 E

Concentration (versus SCE) (versus SCE)

(M) (V) (V)

Thiourea 10-1

0.11 -0.07

10-2

0.11 0.05

Thiocyanate 10-1

0.47 0.30

10-2

0.47 0.42

Thiosulfate 10-1

-0.21 -0.47

10-2

-0.21 -0.27

Chloride 10-1

0.76 0.64

10-2

0.76 0.65

Cyanide 10-1

-0.79 -0.85

10-2

-0.79 -0.73

Combination of the above electrode potential for gold thiosulfate with the reduction potential of

activated carbon (-0.14 versus SCE) gives a negative reduction potential for the gold thiosulfate species

adsorbed onto activated carbon. Thermodynamically, therefore, the gold thiosulfate complex is stable

and will not be reduced during adsorption. Adsorption may be expected to occur via the same

mechanism as the adsorption of the aurocyanide complex onto activated carbon, that is, by ion pairing.

To circumvent the problem of low loading/low affinity for Au(S2O3)23-

, gold can be adsorbed from a

thiosulfate solution by carbon after adding a small amount of cyanide to the system (Lulham and

Lindsay, 1991). The solution is treated with at least a stoichiometric amount of cyanide ions and the

resultant gold cyanide complex is adsorbed on to carbon or a resin. The gold loaded adsorbent is then

subjected to stripping to recover adsorbed gold. It was claimed that selective recovery of gold can be

achieved from thiosulfate feed solution containing other metal ions in solution such as silver, copper

and zinc and 95-97% gold recovery was reported by this method.

Resins

Most anion exchange resins would not be expected to effectively adsorb gold thiosulfate complexes

because thiosulfate is an effective eluant for stripping gold from loaded resins. In general, studies have

found that gold and silver can only be adsorbed on anion exchange at very dilute thiosulfate

concentrations (Wan et al., 1993) and is inhibited by the presence of tetrathionate, a decomposition

product of thiosulfate. (O’Malley, Pers. Comm., PhD student, Murdoch University)

However, Thomas et al (1998) claimed to have achieved almost full recovery of gold and silver using

strong base resins.


39

A strong base resin consisting of a quaternary amine attached to a polymer backbone (e.g. polystyrene

beads) is preferred rather than a weak base resin. This is because weak base resins generally require a

pH of less than 8 and do not have the same capacity as strong resins. In addition, strong bases are more

widely available. Suitable resins include all commerical strong base resins of either type I

(triethylamine functional groups) or type II (triethyl ethanolamine functional groups). To facilitate

screening, macroporous resins with beads of at least 0.8mm in average diameter are preferred. Specific

strong base ion exchange resins for use are listed in Table 7.

Table 7: Suggested resins for gold thiosulfate adsorption (after Thomas et al., 1998)

Resin Type Manufacture

Dowex M-41 Type I Dow chemical

Dowex MSA-1 Type II Dow chemical

Amberlite IRA-904 Type I Rohm&Haas

Amberlite IRA-910 Type II Rohm&Haas

Lewatit M-600, MP 500 Bayer

Gel-type resin 21K Dow chemical

The gold-bearing lixiviant and solid residue are subjected to resin-in-pulp (RIP) or resin-in-leach (RIL)

which recovers both copper and gold.

A schematic diagram of the resin-in-pulp circuit as outlined by Thomas et al (1998) is shown in Figure

17. The RIP operation is carried out in a stirred tank reactor vessel, preferably in a Pachuca tank, which

is an air-agitated, conical bottom, solid-liquid mixing vessel in which the air is injected into the bottom

of the cone. The Pachuca system reduces resin bead breakage and improves dispersion of the resin

beads in the slurry compared with a mechanical agitator system. Between four to eight pachucas are

connected in series with a total residence time for pulp of about 12 hours. Alternatively the resin is

moved from stage to stage counter-currently to the pulp at a rate such that the retention time of the

resin is about two to three hours per stage.

Thomas et al (1998) found that higher gold recoveries were obtained by using very dilute thiosulfate

solutions at 45-55C and adding the resin to the pregnant solution containing the solid residue rather

than adding it to the pregnant solution after a liquid/solid separation step. The presence of the resin in

contact with both the pregnant solution and solid residue reduces degradation of gold thiosulfate and

thiosulfate reagent and reduces the adsorption of gold on the solids.

The pulp and barren lixiviant exiting during the final RIP stage are sent to tailings for further treatment

and reclaim. The loaded resin is sent to an elution stage where copper is eluted from the loaded resin by

washing with ammonia salt or a mixture containing 100-200 g/L of ammonium thiosulfate solution.

The eluate containing ammonium thiosulfate and copper (500-1500 ppm) together with about 10% of

the gold on the resin is then recycled to the leaching circuit.

After copper elution, a thiocyanate solution (100-200g/L) is used to elute gold from the resin. The

eluate is passed to a gold recovery process such as electro-winning whilst the resin is recycled to the

RIP circuit. Because the resin is returned in its SCN- form, it is not clear whether gold is taken up on to

the resin as the Au(SCN)2- complex.


40

Fig 17 Schematic diagram of resin-in-pulp circuit (after Thomas et al., 1998)

ELECTROWINNING

Some work has been carried out on electrowinning gold from thiosulfate solution. The main

electrochemical reaction occurring at the cathode surface is the reduction of the auro-thiosulfate

complex (Gallagher et al., 1989, Abbruzese et al., 1995). The electro reduction of gold thiosulfate to

gold is kinetically faster than the reduction of gold cyanide to gold (Sullivan and Kohl, 1997). Gold

thiosulfate is reduced at -0.15V vs SCE and is independent of pH for values above 4. But because of

the various oxidation and reduction reactions of S2O32-

which occur at the anode and cathode,

electrowinning is not an attractive option for recovery of gold.

CONCLUSIONS

Interest in the use of non-cyanide methods for dissolution of gold is due to the increasing concern

regarding the toxicity of cyanide and the inability of cyanide solution to effectively leach carbonaceous

or complex ores.

Thiosulfate leaching can be considered as a non-toxic alternative to conventional cyanidation. Leaching

by thiosulfate permits a decreasing interference from cations such as lead, zinc, and copper. In some

cases the gold dissolution rates can be faster than for conventional cyanide treatment. The main

disadvantage of thiosulfate however, has been reagent consumption and the lack of a suitable gold

recovery method.

Thiosulfate in the presence of oxygen in solution can leach gold. However without the presence of

ammonia, passivation of gold occurs through the breakdown of thiosulfate to form sulfur coating on the

gold particles. Copper(II) is also required to significantly increase the rate of gold dissolution.

Tails

And

Reclaim

RIP

Au + Cu

Leach

Cu

Elution

Au Elution

Ammonium

Thiosulfate

Potassium

Thiocyanate

Au - Bearing

Eluate to

Au recovery

Au and Cu

Bearing Resin

Resin

Return

Pulp

Cu/(NH4)2S2O3


41

The chemistry of the ammonia-thiosulfate and copper system is complicated due to the simultaneous

presence of complexing ligands such as ammonia and thiosulfate and the Cu(II)-Cu(I) redox couple.

Oxidative decomposition reactions of thiosulfate occur which involve the formation of additional sulfur

compounds such as tetrathionate.

The leaching of gold or silver with ammoniacal thiosulfate can be described as an electrochemical-

catalytic mechanism. Thiosulfate ions react with Au ions formed at anodic sites on the gold surface.

Gold enters the solution to form Au(S2O3)23-

. The Cu(NH3)42+

acquires electrons and is directly reduced

to Cu(NH3)2+ at cathodic sites on the gold surface. In the presence of S2O3

2-, an equilibrium exist

between Cu(NH3)2+ and Cu(S2O3)3

5- ions. Both the Cu(S2O3)3

5- and Cu(NH3)2

+ species in solution are

then regenerated into Cu(NH3)42+

with oxygen.

The mechanism is further complicated by the oxidation of thiosulfate to tetrathionate. Both oxygen and

Cu(NH3)42+

oxidise thiosulfate ions to tetrathionate ions.

Efficient leaching is achieved by maintaining the appropriate concentrations of ammonia and

thiosulfate in solution with copper (II) acting as oxidant. Oxygen is required to maintain the required

Eh to leach gold and convert the reduced cuprous ion to the cupric state for further gold leaching.

Leaching should be carried out at pH values greater than 8.5 to allow free ammonia to be present to

complex copper (II) and reduce the interference of some foreign cations such as iron or manganese. For

effective gold dissolution an oxidation potential between 150-250mV with reference to standard

hydrogen electrode (SHE) is required (depending upon pH). At higher potentials there is significant

oxidation of thiosulfate to tetrathionate.

Higher gold and silver dissolution rates have been reported at temperatures in the range of 40-60C.

Most work has been carried out on complex gold ores containing high copper, carbonaceous ores or

ores containing high concentrations of lead, zinc or manganese. Relatively high concentrations of

thiosulfate, compared with those used in cyanide are generally used, but recent work favours low

concentrations of reagents around 0.1-0.2M thiosulfate and ammonia.

In ammoniacal thiosulfate leaching of gold, increasing the concentration of thiosulfate in solution

results in an increase in thiosulfate consumption. Efforts have been made to reduce consumption by

reducing the concentration of thiosulfate in solution and leaching for extended periods (e.g. 24 hours).

Alternatively, sulfite has been added to the leaching solution to stabilise the thiosulfate.

In the case of low-grade carbonaceous ores, processing by thiosulfate leaching without pretreatment by

heap leaching can be carried out. However, ores containing high sulfide concentrations require at least

partial oxidation prior to thiosulfate leaching. Several investigators have looked at the use of bacteria or

pressure leaching prior to thiosulfate leaching. Another possibility is to generate thiosulfate in-situ. The

stability field for thiosulfate is located in the neutral to basic pH region. Once thiosulfate has been

formed conditions can be readily adjusted for gold and silver leaching. In addition to thiosulfate,

polysulfides will also be present during leaching which provide another complexant for leaching gold

and silver.

While thermodynamic predictions can be made, research is required on detailing the reactions and

species involved in thiosulfate leaching of gold and silver in order to establish the optimum molar

ratios for ammonia, thiosulfate and copper. In addition, the oxygen concentration required in the

ammonical thiosulfate and copper systems to regenerate copper to the Cu(II) state requires

investigation.

Most research has been concerned with the extractability of gold with thiosulfate and only limited work

has been carried out on the process of recovering gold from solutions. Precipitation methods appear to

have been used the most and are relatively successful on clarified liquors. Alternative processes include

a reduction-precipitation process with stabilised alkali metal borohydrides, hydrogen sparging, use of

alkyl phosphorus esters, extraction using amines, addition of sulfide to solution, and electro winning to

remove gold from solution. However in most cases coprecipitation of copper occurs which gives a low

grade gold product or high reagent consumption. Activated carbon has a low affinity for Au(S2O3)23-

.


42

However, gold complex anions can be adsorbed from a thiosulfate solution by carbon after adding a

small amount of cyanide to the system.

The use of a strong base resin consisting of a quaternary amine attached to a polymer backbone has

been partially successful in resin-in-pulp studies. Unfortunately copper also loads on to the resin and

thiosulfate ion inhibits the loading of gold. More research into the selective adsorption or stripping of

gold is required.

The difficulties described here do not appear insurmountable given a detailed understanding of the

mechanism involved which should lead to an effective alternative and less toxic methodology for gold

extraction.

ACKNOWLEDGEMENTS

Advice on the construction of speciation diagrams by Dr Ewen Silvester is gratefully appreciated.


43

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thiosulfate leaching of gold – a review

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